CN102747226B - Method for treating zinc hydrometallurgy waste residue by using alkali ammonium sulfur coupling method - Google Patents
Method for treating zinc hydrometallurgy waste residue by using alkali ammonium sulfur coupling method Download PDFInfo
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- CN102747226B CN102747226B CN201210122449.6A CN201210122449A CN102747226B CN 102747226 B CN102747226 B CN 102747226B CN 201210122449 A CN201210122449 A CN 201210122449A CN 102747226 B CN102747226 B CN 102747226B
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- zinc
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- ammonium
- ammonium chloride
- waste residue
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- 238000000034 method Methods 0.000 title claims abstract description 98
- 239000011701 zinc Substances 0.000 title claims abstract description 75
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 71
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 68
- 239000003513 alkali Substances 0.000 title claims abstract description 64
- 239000002699 waste material Substances 0.000 title claims abstract description 42
- 238000009854 hydrometallurgy Methods 0.000 title claims abstract description 36
- HIVLDXAAFGCOFU-UHFFFAOYSA-N ammonium hydrosulfide Chemical compound [NH4+].[SH-] HIVLDXAAFGCOFU-UHFFFAOYSA-N 0.000 title claims abstract description 22
- 238000010168 coupling process Methods 0.000 title abstract description 4
- NLXLAEXVIDQMFP-UHFFFAOYSA-N Ammonia chloride Chemical compound [NH4+].[Cl-] NLXLAEXVIDQMFP-UHFFFAOYSA-N 0.000 claims abstract description 141
- 235000019270 ammonium chloride Nutrition 0.000 claims abstract description 74
- 239000007788 liquid Substances 0.000 claims abstract description 55
- 239000010949 copper Substances 0.000 claims abstract description 37
- 229910052709 silver Inorganic materials 0.000 claims abstract description 36
- 229910052802 copper Inorganic materials 0.000 claims abstract description 35
- 239000004332 silver Substances 0.000 claims abstract description 32
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 30
- 229910052793 cadmium Inorganic materials 0.000 claims abstract description 29
- 239000011133 lead Substances 0.000 claims abstract description 27
- BDOSMKKIYDKNTQ-UHFFFAOYSA-N cadmium atom Chemical compound [Cd] BDOSMKKIYDKNTQ-UHFFFAOYSA-N 0.000 claims abstract description 25
- 238000001556 precipitation Methods 0.000 claims abstract description 23
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 19
- 239000012141 concentrate Substances 0.000 claims abstract description 11
- 239000005083 Zinc sulfide Substances 0.000 claims abstract description 8
- DRDVZXDWVBGGMH-UHFFFAOYSA-N zinc;sulfide Chemical compound [S-2].[Zn+2] DRDVZXDWVBGGMH-UHFFFAOYSA-N 0.000 claims abstract description 7
- 239000002893 slag Substances 0.000 claims description 80
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 48
- 230000008569 process Effects 0.000 claims description 42
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 claims description 30
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 24
- 238000007654 immersion Methods 0.000 claims description 24
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 claims description 23
- 241000370738 Chlorion Species 0.000 claims description 18
- 238000000926 separation method Methods 0.000 claims description 18
- 238000002156 mixing Methods 0.000 claims description 15
- 235000011114 ammonium hydroxide Nutrition 0.000 claims description 13
- 229910052742 iron Inorganic materials 0.000 claims description 13
- 239000000203 mixture Substances 0.000 claims description 13
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 claims description 12
- 238000001914 filtration Methods 0.000 claims description 12
- 150000002500 ions Chemical class 0.000 claims description 12
- UQSXHKLRYXJYBZ-UHFFFAOYSA-N iron oxide Inorganic materials [Fe]=O UQSXHKLRYXJYBZ-UHFFFAOYSA-N 0.000 claims description 11
- 229910000359 iron(II) sulfate Inorganic materials 0.000 claims description 11
- SURQXAFEQWPFPV-UHFFFAOYSA-L iron(2+) sulfate heptahydrate Chemical compound O.O.O.O.O.O.O.[Fe+2].[O-]S([O-])(=O)=O SURQXAFEQWPFPV-UHFFFAOYSA-L 0.000 claims description 10
- GEHJYWRUCIMESM-UHFFFAOYSA-L sodium sulfite Chemical compound [Na+].[Na+].[O-]S([O-])=O GEHJYWRUCIMESM-UHFFFAOYSA-L 0.000 claims description 10
- NDLPOXTZKUMGOV-UHFFFAOYSA-N oxo(oxoferriooxy)iron hydrate Chemical compound O.O=[Fe]O[Fe]=O NDLPOXTZKUMGOV-UHFFFAOYSA-N 0.000 claims description 9
- 229910001308 Zinc ferrite Inorganic materials 0.000 claims description 7
- PIJPYDMVFNTHIP-UHFFFAOYSA-L lead sulfate Chemical compound [PbH4+2].[O-]S([O-])(=O)=O PIJPYDMVFNTHIP-UHFFFAOYSA-L 0.000 claims description 7
- 229910052981 lead sulfide Inorganic materials 0.000 claims description 7
- 229940056932 lead sulfide Drugs 0.000 claims description 7
- 238000004064 recycling Methods 0.000 claims description 7
- WGEATSXPYVGFCC-UHFFFAOYSA-N zinc ferrite Chemical compound O=[Zn].O=[Fe]O[Fe]=O WGEATSXPYVGFCC-UHFFFAOYSA-N 0.000 claims description 7
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 claims description 7
- 229960001763 zinc sulfate Drugs 0.000 claims description 7
- 229910000368 zinc sulfate Inorganic materials 0.000 claims description 7
- 229910021529 ammonia Inorganic materials 0.000 claims description 6
- UYJXRRSPUVSSMN-UHFFFAOYSA-P ammonium sulfide Chemical group [NH4+].[NH4+].[S-2] UYJXRRSPUVSSMN-UHFFFAOYSA-P 0.000 claims description 5
- 235000010265 sodium sulphite Nutrition 0.000 claims description 5
- 239000012535 impurity Substances 0.000 claims description 4
- 230000002829 reductive effect Effects 0.000 claims description 2
- 238000002386 leaching Methods 0.000 abstract description 15
- 238000011084 recovery Methods 0.000 abstract description 15
- 229910052751 metal Inorganic materials 0.000 abstract description 14
- 239000002184 metal Substances 0.000 abstract description 14
- 230000008901 benefit Effects 0.000 abstract description 6
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 abstract description 5
- 230000007613 environmental effect Effects 0.000 abstract description 4
- LWUVWAREOOAHDW-UHFFFAOYSA-N lead silver Chemical group [Ag].[Pb] LWUVWAREOOAHDW-UHFFFAOYSA-N 0.000 abstract description 3
- 239000007787 solid Substances 0.000 abstract description 3
- 230000008878 coupling Effects 0.000 abstract description 2
- 238000005859 coupling reaction Methods 0.000 abstract description 2
- 229910052984 zinc sulfide Inorganic materials 0.000 abstract 2
- LCPUDZUWZDSKMX-UHFFFAOYSA-K azane;hydrogen sulfate;iron(3+);sulfate;dodecahydrate Chemical group [NH4+].O.O.O.O.O.O.O.O.O.O.O.O.[Fe+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O LCPUDZUWZDSKMX-UHFFFAOYSA-K 0.000 abstract 1
- 238000001354 calcination Methods 0.000 abstract 1
- 231100001261 hazardous Toxicity 0.000 abstract 1
- JQJCSZOEVBFDKO-UHFFFAOYSA-N lead zinc Chemical compound [Zn].[Pb] JQJCSZOEVBFDKO-UHFFFAOYSA-N 0.000 abstract 1
- 230000001376 precipitating effect Effects 0.000 abstract 1
- 229910052745 lead Inorganic materials 0.000 description 6
- 239000012071 phase Substances 0.000 description 6
- 229910017840 NH 3 Inorganic materials 0.000 description 5
- 238000006243 chemical reaction Methods 0.000 description 5
- 238000012545 processing Methods 0.000 description 5
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 4
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 4
- 229910052785 arsenic Inorganic materials 0.000 description 3
- 238000000151 deposition Methods 0.000 description 3
- 238000005188 flotation Methods 0.000 description 3
- 229910001385 heavy metal Inorganic materials 0.000 description 3
- 239000007791 liquid phase Substances 0.000 description 3
- 238000011160 research Methods 0.000 description 3
- 239000000779 smoke Substances 0.000 description 3
- 231100000331 toxic Toxicity 0.000 description 3
- 230000002588 toxic effect Effects 0.000 description 3
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 3
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 2
- 239000011449 brick Substances 0.000 description 2
- 239000004568 cement Substances 0.000 description 2
- YCKOAAUKSGOOJH-UHFFFAOYSA-N copper silver Chemical compound [Cu].[Ag].[Ag] YCKOAAUKSGOOJH-UHFFFAOYSA-N 0.000 description 2
- 230000008021 deposition Effects 0.000 description 2
- 238000005265 energy consumption Methods 0.000 description 2
- 239000002241 glass-ceramic Substances 0.000 description 2
- 229910052738 indium Inorganic materials 0.000 description 2
- 229910052935 jarosite Inorganic materials 0.000 description 2
- 238000004519 manufacturing process Methods 0.000 description 2
- 238000005554 pickling Methods 0.000 description 2
- 231100000614 poison Toxicity 0.000 description 2
- 239000000377 silicon dioxide Substances 0.000 description 2
- 229960001866 silicon dioxide Drugs 0.000 description 2
- 235000012239 silicon dioxide Nutrition 0.000 description 2
- 239000010944 silver (metal) Substances 0.000 description 2
- 239000007790 solid phase Substances 0.000 description 2
- 238000007711 solidification Methods 0.000 description 2
- 230000008023 solidification Effects 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 description 1
- 208000035126 Facies Diseases 0.000 description 1
- FOIXSVOLVBLSDH-UHFFFAOYSA-N Silver ion Chemical compound [Ag+] FOIXSVOLVBLSDH-UHFFFAOYSA-N 0.000 description 1
- 238000003723 Smelting Methods 0.000 description 1
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical group [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- 238000009825 accumulation Methods 0.000 description 1
- 229910052787 antimony Inorganic materials 0.000 description 1
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 description 1
- APAWRDGVSNYWSL-UHFFFAOYSA-N arsenic cadmium Chemical compound [As].[Cd] APAWRDGVSNYWSL-UHFFFAOYSA-N 0.000 description 1
- 238000011109 contamination Methods 0.000 description 1
- 239000013078 crystal Substances 0.000 description 1
- 238000000354 decomposition reaction Methods 0.000 description 1
- 238000011161 development Methods 0.000 description 1
- 230000000694 effects Effects 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 238000001704 evaporation Methods 0.000 description 1
- 230000008020 evaporation Effects 0.000 description 1
- 238000010304 firing Methods 0.000 description 1
- 239000003500 flue dust Substances 0.000 description 1
- 239000003546 flue gas Substances 0.000 description 1
- 238000011835 investigation Methods 0.000 description 1
- RUTXIHLAWFEWGM-UHFFFAOYSA-H iron(3+) sulfate Chemical compound [Fe+3].[Fe+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O RUTXIHLAWFEWGM-UHFFFAOYSA-H 0.000 description 1
- 229910000360 iron(III) sulfate Inorganic materials 0.000 description 1
- 238000007885 magnetic separation Methods 0.000 description 1
- 229910052976 metal sulfide Inorganic materials 0.000 description 1
- 238000005272 metallurgy Methods 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- 239000012452 mother liquor Substances 0.000 description 1
- 238000009856 non-ferrous metallurgy Methods 0.000 description 1
- 150000002927 oxygen compounds Chemical class 0.000 description 1
- 239000000049 pigment Substances 0.000 description 1
- 239000002574 poison Substances 0.000 description 1
- 230000007096 poisonous effect Effects 0.000 description 1
- 239000000843 powder Substances 0.000 description 1
- 239000010970 precious metal Substances 0.000 description 1
- 238000002360 preparation method Methods 0.000 description 1
- 229940083025 silver preparation Drugs 0.000 description 1
- 238000005245 sintering Methods 0.000 description 1
- 239000002910 solid waste Substances 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 229910052717 sulfur Inorganic materials 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- 238000005987 sulfurization reaction Methods 0.000 description 1
- 125000000101 thioether group Chemical group 0.000 description 1
- 231100000167 toxic agent Toxicity 0.000 description 1
- 239000003440 toxic substance Substances 0.000 description 1
- 239000001052 yellow pigment Substances 0.000 description 1
Classifications
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Abstract
The present invention discloses a method for treating zinc hydrometallurgy waste residue by using an alkali ammonium sulfur coupling method, relates to a method for comprehensive utilization of lead silver residue and iron alum residue in zinc hydrometallurgy waste residue by using a alkali ammonium sulfur coupling clean metallurgical method, and belongs to the field of metallurgical hazardous solid reside treatment, wherein the method can further be used to treat other complex waste containing lead and zinc. According to the method, a medium temperature calcination method is performed on zinc hydrometallurgy residue; then an alkali leaching treatment is performed; then the leached alkali residue is leached with ammonium chloride; the resulting alkali leaching liquid is replaced by zinc powder to obtain residues containing zinc, copper, lead, silver and cadmium, and a solution; the solution is subjected to precipitation by using a sulfide to obtain a zinc sulfide product and an ammonium chloride solution; or after the ammonium chloride is leached, the resulting leaching liquid is directly treated by using a precipitating agent, and then is filtered to obtain a lead-zinc sulfide concentrate containing copper and silver and an ammonium chloride solution. With the method of the present invention, valuable metal resources in zinc hydrometallurgy residue can be subjected to effective comprehensive recovery, harmlessness and resource of the final residue are achieved, and good environmental benefits and economic benefits can be achieved.
Description
Technical field
The invention belongs to metallurgical dangerous solid slag process field, the particularly plumbous silver-colored slag of clean metallurgical method comprehensive utilization zinc hydrometallurgy waste residue of a kind of alkali ammonium sulphur coupling and the method for iron vitriol slag, also can be used for processing the complicated waste material of the leaded zinc of other type.
Background technology
Plumbous silver-colored slag and iron vitriol slag are two kinds of solid slags of output in Zinc Hydrometallurgy Process, owing to wherein not only containing the heavy metal ion such as Zn, Cu, but also contain the poisonous and harmful elements such as Pb, Cd and As, and environmental stability is poor, part Toxic has water-soluble mutually, therefore, classified in the world as dangerous solid waste.According to coloured Industrial Development Bureau statistics, the zinc annual production of China in 2009 surpasses 4,400,000 tons, occupies first place in the world.Wherein, the zinc that wet process is produced accounts for greatly 75% of total amount, will discharge every year above-mentioned two kinds of waste residues and reach more than 2,000,000 tons.But the treatment process of these slags also exists economically or technical problem at present, except some are containing In, the more lead silver slag of the rare precious metals such as Ge adopts outside rotary kiln fuming process simple process, approximately has more than 60% a large amount of waste residues all to pile and have smeltery's periphery for a long time every year.
At present, mainly contain four kinds of industrial application for the treatment process of this class waste residue, have been realized: take fuming process, flotation process and the solidification method that two sections of Ausmelt oven processes, China that the Korea S Osan of Xin Ye company smeltery is representative generally adopt.
Two sections of Ausmelt oven process process metal rate of recovery are higher, wherein, and zinc recovery 86%, lead recovery 91%, silver raising recovery rate 88%, also recyclable part copper, antimony etc.But its shortcoming is that investment in fixed assets is large, smelting temperature is high, the smoke pollution problem of sulfur-bearing arsenic cadmium and the scum grade of output low, can only be for cement industry etc.
China zinc hydrometallurgy enterprise adopts rotary kiln fuming process more, with reductive agents such as a large amount of broken coals, under the high temperature of 1100～1300 ℃, the element evaporations such as Zn, Pb, Cd and In in plumbous silver-colored slag is entered to flue dust, then processes.But the method energy consumption is large, and smoke pollution is large, also can cause the waste of the resources such as part Cu, Ag.
Zhu Zhou smeltery once adopted flotation process to reclaim the silver in plumbous silver-colored slag, can obtain the silver preparation concentrate containing Ag 0.6% left and right, silver raising recovery rate can reach 76.5% left and right, but in slag, other valuable metal all can not get efficient recovery as Zn, Pb and Cu etc., flotation dreg is not detoxified yet, and also needs further processing.
It is simple and easy to do that cement solidification method is processed zinc hydrometallurgy waste residue, in some factory of China, adopted; In India and Italy, investigator adopts roasting-sintering process to be made into siderotil brick or glass-ceramic that intensity is higher.But these treatment processs are all failed valuable metal in efficient recovery slag, and can emit toxic oxygen compound in high-temperature firing process, and the potential risk that discharges toxic substance in environment after existing siderotil brick, glass-ceramic to deposit for a long time.
Technical process from industrialization at present, mainly with thermal process, be treated to master, subject matter is that flue gas, smoke contamination are more serious, and China domestic enterprise thermal process also exists energy consumption high, the problems such as the comprehensive recovery of valuable metal is not high, and economy is poor.
For the resource better, in more clean this class danger wastes of comprehensive utilization, Chinese scholars also carrying out a large amount of deep research, has proposed a series for the treatment of processs and technique up to now.Be summed up and can be divided into pickling process, alkaline leaching and chlorinating roasting etc.
Pickling process mainly contains following a few class:
Sulfuric acid leaching-soak the slag technical process that in sulfuric acid prepared by microwave radiation-iron powder slaking-bodied ferric sulfate to process iron vitriol slag; The method of the high Ore Leaching-iron replacement of high temperature-copperas solution precipitation is processed the technique that iron vitriol slag is produced iron oxide yellow pigments.These two kinds of techniques make the iron resources in iron vitriol slag obtain effective utilization.But in these techniques slag containing the recovery of valuable metal, iron replacement slag how economical and effective process and the aspects such as economy of technique need further investigation.
Also have scholar to propose first alkali and soak, then selectivity acidleach Zn and In, last scum magnetic separation obtains the full wet processing of iron ore concentrate.Valuable metal in not only recyclable iron vitriol slag but also can make the iron resources in slag be utilized preferably.But also need the reuse of alkali immersion liquid and how to reclaim valuable metal in alkali immersion liquid etc. and do further research.
By research trends, the advantage such as Wet-process metallurgy method is processed these waste residues owing to having environmental facies to close friend, and resource recovery is high, fixed investment is few, and become the main direction of studying of current processing iron vitriol slag and plumbous silver-colored slag.
In sum, how the comprehensive from zinc hydrometallurgy waste residue of economical and efficient reclaimed various valuable metals, make tailings without the study hotspot and the difficult point that poison resource utilization even and remain current Non-ferrous Metallurgy field simultaneously, there is very important theory significance and realistic meaning.
summary of the invention:
The present invention proposes alkali ammonium sulphur coupled method and process the method for zinc hydrometallurgy waste residue, particularly alkali ammonium sulphur coupled method is processed the technical process of iron vitriol slag and plumbous silver-colored slag, can solve the problem of Fe and the separated of valuable metal and the recycle of medium system in these waste residue processes of wet processing.
Technical scheme of the present invention is: it is ZnSO that middle low-temperature bake makes the principal elements such as Zn in waste residue, Pb and Fe compose the thing phase decomposition of depositing
4, PbSO
4and Fe
2o
3; After sig water leaches roasting, slag is to remove As wherein
3+, K
+, SO
4 2-and Mg
2+, avoid these ions at follow-up NH
4accumulation in Cl system; NH
4cl leaching process is to utilize Cl
-with Ag, Pb, NH
3with the strong coordination of Zn, Cu, Ag and Cd, and make these valuable metal strippings in waste residue in liquid phase; In leach liquor, the recovery of valuable metal has two kinds of methods: (1) direct sulfide precipitation generates the lead sulfide zinc ore concentrate of rich silver-bearing copper; (2) first use Zn powder substitution Pb wherein
2+, Cu
2+, Ag
+and Cd
2+, and then with sulfide by NH
4zn in Cl solution
2+precipitated crystal is pigment-level ZnS product, and mother liquor of precipitation of ammonium is mainly containing NH simultaneously
4cl can turn back to ammonium chloride and leach operation; Remove after As sig water adopts the dearsenification of sulfuration method to process after repeatedly circulating and return to again use; Finishing slag is mainly convenient to dispose and recycling containing ferric oxide and silicon-dioxide.
Alkali ammonium sulphur coupled method of the present invention is processed the method for zinc hydrometallurgy waste residue, and concrete steps comprise as follows:
(1) by the roasting 0.5～3 hour at 300～900 ℃ of temperature of the waste residue of zinc hydrometallurgy, it is the fired slags of zinc sulfate, lead sulfate and ferric oxide that middle low-temperature bake makes siderotil or zinc ferrite thing inversion of phases in slag;
(2) by solid-to-liquid ratio (g/ml) 1:4～1:10, with sodium hydroxide solution, at 30～160 ℃, leach the fired slags that obtains in step (1) 0.5～3 hour, sig water leaching fired slags is to remove As wherein
3+, K
+, SO
4 2-and Mg
2+, solid-liquid separation obtains alkali and soaks slag and alkali immersion liquid, and alkali immersion liquid can be returned to recycled for multiple times (repeatedly can adopt prior art dearsenification after circulation); The reaction wherein occurring is:
As
2O
3?+?NaOH?→?2NaAsO
2(aq)?+?H
2O
(3) alkali step (2) being obtained soaks slag and immerses 0.5～3h in the mixing solutions that temperature is 90～125 ℃ of ammonium chlorides and ammoniacal liquor, the leached mud that obtains containing the leach liquor of zinc, copper, lead, cadmium and silver etc. after filtration and mainly contain ferric oxide and silicon-dioxide, in slag, not containing toxic heavy metal, can be used as the batching of iron ore concentrate; Its chemical reaction is:
Me
x+?+?m?NH
4Cl?=?Me(NH
3)
i(Cl)
j x-j?+?(m-i)NH
3?+?(m-j)Cl
-?+?m?H
+
Me wherein
x+represent Zn
2+, Pb
2+, Cu
2+, Ag
+and Cd
2+; X is the ionic valence condition of Me; I=0,1,2,3,4,5 or 6, it represents the NH with Me coordination
3quantity; J=0,1,2,3,4,5 or 6, it represents the Cl with Me coordination
-quantity, m is i and the large number of j intermediate value.
(4) according to the ionic concn of copper, lead, cadmium and silver in ammonium chloride leach liquor, be scaled total mol concentration, with the zinc powder that mole number is 0.5～3.0 times of above-mentioned total mole number, at 20～60 ℃, reduce 5～60 minutes to reclaim copper, lead, cadmium and the silver in leach liquor, after solid-liquid separation, obtain respectively the cadmia of leaded, copper, cadmium and silver and containing the ammonium chloride solution of zine ion;
Principal reaction is: Me (NH
3)
i(Cl)
j x-j+ Zn=Zn (NH
3)
i(Cl)
j x-j+ Me
Wherein Me represents Pb, Cu, Ag and Cd, the ionic valence condition that x is Me, and i represents the NH with Me coordination
3quantity, j represents the Cl with Me coordination
-quantity.
(5) by ammonium chloride solution in step (4), the precipitation agent containing 0.5～2.0 times of zine ion mole number adds in ammonium chloride solution, react and be precipitated thing after 5～30 minutes, solid-liquid separation, obtaining respectively purity is more than 99% zinc sulphide product and the ammonium chloride leach liquor can be recycled.(as shown in Figure 1)
Principal reaction is: Me
2++ S
2-=MeS ↓.
The ammonium chloride leach liquor (as shown in Figure 2) obtaining in step (3), can also adopt precipitation agent directly to process 5～30 minutes, precipitation agent consumption is zinc, lead, copper, cadmium and silver-colored integral molar quantity 0.5～1.5 times in ammonium chloride, (zinc content is 20 ~ 60% after filtration, to obtain being rich in the lead sulfide zinc ore concentrate of copper silver, content 100 ~ 500 g/t of silver, content 0.3 ~ 3.0 wt% of copper) and ammonium chloride solution.Principal reaction is: Zn
2++ S
2-=ZnS ↓
The waste residue of described zinc hydrometallurgy is plumbous silver-colored slag or iron vitriol slag, its composition and percentage range: zinc 3 ~ 15%, and plumbous 3～10 wt%, silver 50～500 g/t, iron 5～30 wt%, all the other are impurity.
The concentration of described sodium hydroxide solution is 30～100 g/L.
In described step (2), obtain alkali immersion liquid recycle time, concentration is during lower than 30 g/L, recycling while adding sodium hydroxide to adjust concentration to 30～100 g/L.
In described step (3), ammonia concn is 1
~30g/L, the concentration of ammonium chloride solution is 100～400g/L, in mixing solutions, the mol ratio of ammonium ion and chlorion is 1:1～2:1.
In described step (3), alkali soaks in slag immersion ammonium chloride and ammoniacal liquor mixing solutions, and solid-to-liquid ratio is 1:10(g/ml).
The granularity of described zinc powder is-100 orders, and purity is more than 99%.
Described precipitation agent is sulfide: ammonium sulfide or sodium sulphite, be common commercially available.
The ammonium chloride solution that described two kinds of separation methods finally obtain all can return to step (2) and recycle, and when the mol ratio of ammonium ion and chlorion is during lower than 1:1, the mol ratio that adds ammoniacal liquor to be adjusted to ammonium ion and chlorion is: 2:1～1:1; When mol ratio is during higher than 2:1, the mol ratio that adds ammonium chloride to be adjusted to ammonium ion and chlorion is: 2:1～1:1.
Advantage of the present invention and positively effect:
The novel process that the present invention proposes, can comprehensively reclaim the valuable metal resource in zinc hydrometallurgy waste residue effectively, and heavy metal and the rate of recovery can reach more than 95%, make finishing slag innoxious simultaneously, and resource utilization, can realize good environmental benefit and economic benefit.
Accompanying drawing explanation
Fig. 1 is that alkali ammonium sulphur coupled method of the present invention is processed the principle process flow sheet that zinc hydrometallurgy waste residue is prepared zinc sulphide product.
Fig. 2 is that alkali ammonium sulphur coupled method of the present invention is processed the principle process flow sheet that the lead sulfide zinc ore concentrate of copper, silver is rich in the preparation of zinc hydrometallurgy waste residue.
Embodiment
Below in conjunction with embodiment and accompanying drawing, the present invention will be further described, but the invention is not restricted to the following stated scope.
Embodiment 1:
The alkali ammonium sulphur coupled method of the present embodiment is processed the method for zinc hydrometallurgy waste residue, has adopted the Jarosite Residues of output in the Lead And Zinc Smelter production process of northwest, and its chemical composition is as shown in table 1:
(1) by Jarosite Residues roasting 1 hour at 650 ℃ of temperature, making siderotil or zinc ferrite thing inversion of phases in slag is the fired slags of zinc sulfate, lead sulfate and ferric oxide, its burn after composition as shown in the first row in table 2.
Note: in table, " ^ " representation unit is that g/t (solid phase) or ppm (liquid phase) "-" represent that the amount detecting is extremely low, and " * " represents to detect (is mainly the SO in solution
4 2-content is not surveyed).
(2) by solid-to-liquid ratio 1:4, by concentration, be that the sodium hydroxide solution of 50 g/L leaches the fired slags that obtains in step (1) 1 hour at 100 ℃, solid-liquid separation obtains alkali and soaks slag and alkali immersion liquid, and recycle is returned in alkali immersion liquid; The composition of gained leached mud is as shown in table 2 the second row.Slag rate is 85%, and the leaching yield of arsenic reaches 72%.
During alkali immersion liquid recycle, concentration is during lower than 50 g/L, recycling while adding sodium hydroxide to adjust concentration to 50 g/L.
(3) alkali step (2) being obtained soaks slag and immerses 2h in the mixing solutions that temperature is 100 ℃ of ammonium chlorides and ammoniacal liquor, and solid-to-liquid ratio 1:10, obtains ammonium chloride leach liquor and leached mud after filtration; (ammonia concn is 3g/L, and the concentration of ammonium chloride solution is 280g/L, and in mixing solutions, the mol ratio of ammonium ion and chlorion is 1:1.) gained ammonium soak slag composition as shown in table 2 the third line, slag rate is 75%.The composition of leach liquor is as shown in table 2 fourth line.Ammonium soaks zinc in process, copper, plumbous silver and cadmium leaching yield and all reaches more than 97%, and Fe is not leached.Ammonium chloride solution after leaching all can return to step (2) and recycle, and when the mol ratio of ammonium ion and chlorion is during lower than 1:1, the mol ratio that adds ammoniacal liquor to be adjusted to ammonium ion and chlorion is 1:1.
(4) by the zinc powder of 2 times of the total mole numbers of copper, lead, cadmium and silver in ammonium chloride leach liquor, (granularity is-100 orders, purity is more than 99%) at 40 ℃, reduce 40 minutes to reclaim copper, lead, cadmium and the silver in leach liquor, after solid-liquid separation, obtain respectively the cadmia of leaded, copper, cadmium and silver and containing the ammonium chloride solution of zine ion; After reduction, the composition of liquid is as shown in table 2 fifth line, and wherein copper reduction rate can reach 95%, silver raising recovery rate 85%, the rate of recovery 87% of cadmium, plumbous reduction ratio reach 99%.
(5) by ammonium chloride solution in step (4), the precipitation agent (ammonium sulfide) containing 1 times of zine ion mole number adds in ammonium chloride solution, reacts and is precipitated thing after 15 minutes, solid-liquid separation, the zinc sulphide obtaining respectively and ammonium chloride solution; Zn deposition rate reaches 96%.The content of sulfide precipitation slag and solution is as shown in table 2 the 7th and eight row.
Embodiment 2:
The alkali ammonium sulphur coupled method of the present embodiment is processed the method for zinc hydrometallurgy waste residue, adopts the lead silver slag from northwest zinc hydrometallurgy factory, and its chemical composition is as shown in table 3:
(1) by the roasting 1.5 hours at 550 ℃ of temperature of the silver-colored slag of lead, making siderotil or zinc ferrite thing inversion of phases in slag is the fired slags of zinc sulfate, lead sulfate and ferric oxide; It burns rear composition as shown in the first row in table 4.
Note: in table, " ^ " representation unit is that g/t (solid phase) or ppm (liquid phase) "-" represent that the amount detecting is extremely low, and " * " represents to detect (is mainly the SO in solution
4 2-content is not surveyed).
(2) by solid-to-liquid ratio 1:8, by concentration, be that the sodium hydroxide solution of 30 g/L leaches the fired slags that obtains in step (1) 1.2 hours at 160 ℃, solid-liquid separation obtains alkali and soaks slag and alkali immersion liquid, and recycle is returned in alkali immersion liquid; The composition of gained leached mud is as shown in table 4 the second row.Slag rate is 83%, and the leaching yield of arsenic reaches 74%.During alkali immersion liquid recycle, concentration is during lower than 30 g/L, recycling while adding sodium hydroxide to adjust concentration to 30 g/L.
(3) alkali step (2) being obtained soaks slag and immerses 3h in the mixing solutions that temperature is 105 ℃ of ammonium chlorides and ammoniacal liquor, and solid-to-liquid ratio 1:10, obtains ammonium chloride leach liquor and leached mud after filtration; (ammonia concn is 3g/L, and the concentration of ammonium chloride solution is 280g/L, and in mixing solutions, the mol ratio of ammonium ion and chlorion is 1.5:1.) gained ammonium soak slag composition as shown in table 4 the third line, slag rate is 77%.The composition of leach liquor is as shown in table 4 fourth line.Ammonium soaks in process that zinc leaching rate reaches 98%, copper leaching rate 94%, plumbous leaching yield 97%, silver and cadmium leaching yield reach 100%, Fe and do not leached.Ammonium chloride solution after leaching all can return to step (2) and recycle, and when the mol ratio of ammonium ion and chlorion is during lower than 1.5:1, the mol ratio that adds ammoniacal liquor to be adjusted to ammonium ion and chlorion is: 1.5:1.
The ammonium chloride leach liquor (as shown in Figure 2) obtaining in step (3), can adopt precipitation agent (sodium sulphite) directly to process 15 minutes, precipitation agent consumption is zinc, lead, copper, cadmium and silver-colored integral molar quantity 1 times in ammonium chloride, obtains being rich in lead sulfide zinc ore concentrate and the ammonium chloride solution of copper silver after filtration.Wherein the deposition rate of valuable metal is all over 97%.The content of sulfide precipitation slag and solution is as shown in table 4 the 5th and six row.
Embodiment 3:
(1) by the waste residue of zinc hydrometallurgy (silver-colored 50 g/t, iron 30 wt%, all the other are impurity for zinc 3%, plumbous 10 wt%) roasting 3 hours at 300 ℃ of temperature, making siderotil or zinc ferrite thing inversion of phases in slag is the fired slags of zinc sulfate, lead sulfate and ferric oxide;
(2) by solid-to-liquid ratio 1:10, by concentration, be that 100 g/L sodium hydroxide solutions leach the fired slags that obtains in step (1) 0.5 hour at 30 ℃, solid-liquid separation obtains alkali and soaks slag and alkali immersion liquid, and recycle is returned in alkali immersion liquid; During alkali immersion liquid recycle, concentration is during lower than 100 g/L, recycling while adding sodium hydroxide to adjust concentration to 100g/L.
(3) alkali step (2) being obtained soaks slag, and according to solid-to-liquid ratio 1:10, to immerse in the mixing solutions that temperature is 90 ℃ of ammonium chlorides (concentration is 100g/L) and ammoniacal liquor (ammonia concn is 1g/L) mol ratio of ammonium ion and chlorion in 3h(mixing solutions be 2:1), after filtration, obtain ammonium chloride leach liquor and leached mud;
(4) by the zinc powder of 0.5 times of the total mole number of copper, lead, cadmium and silver in ammonium chloride leach liquor, (granularity is-100 orders, purity is more than 99%) at 20 ℃, reduce 60 minutes to reclaim copper, lead, cadmium and the silver in leach liquor, after solid-liquid separation, obtain respectively the cadmia of leaded, copper, cadmium and silver and containing the ammonium chloride solution of zine ion;
(5) by ammonium chloride solution in step (4), the precipitation agent (ammonium sulfide) containing 0.5 times of zine ion mole number adds in ammonium chloride solution, reacts and is precipitated thing after 30 minutes, solid-liquid separation, the zinc sulphide obtaining respectively and ammonium chloride solution; The ammonium chloride solution finally obtaining all can return to step (2) and recycle, and when mol ratio is during higher than 2:1, the mol ratio that adds ammonium chloride to be adjusted to ammonium ion and chlorion is: 2:1.
Embodiment 4:
(1) by the waste residue of zinc hydrometallurgy (silver-colored 500 g/t, iron 5 wt%, all the other are impurity for zinc 15%, plumbous 3 wt%) roasting 0.5 hour at 900 ℃ of temperature, making siderotil or zinc ferrite thing inversion of phases in slag is the fired slags of zinc sulfate, lead sulfate and ferric oxide;
(2) by solid-to-liquid ratio 1:6, by concentration, be that 100 g/L sodium hydroxide solutions leach the fired slags that obtains in step (1) 3 hours at 160 ℃, solid-liquid separation obtains alkali and soaks slag and alkali immersion liquid, and recycle is returned in alkali immersion liquid; During alkali immersion liquid recycle, concentration is during lower than 100 g/L, recycling while adding sodium hydroxide to adjust concentration to 100 g/L.
(3) alkali step (2) being obtained soaks slag, and according to solid-to-liquid ratio 1:10, to immerse in the mixing solutions that temperature is 125 ℃ of ammonium chlorides (concentration is 400g/L) and ammoniacal liquor (ammonia concn is 30g/L) mol ratio of ammonium ion and chlorion in 0.5h(mixing solutions be 2:1), after filtration, obtain ammonium chloride leach liquor and leached mud;
(4) by the zinc powder of 3.0 times of the total mole numbers of copper, lead, cadmium and silver in ammonium chloride leach liquor, (granularity is-100 orders, purity is more than 99%) at 60 ℃, reduce 5 minutes to reclaim copper, lead, cadmium and the silver in leach liquor, after solid-liquid separation, obtain respectively the cadmia of leaded, copper, cadmium and silver and containing the ammonium chloride solution of zine ion;
(5) by ammonium chloride solution in step (4), the precipitation agent (sodium sulphite) containing 2.0 times of zine ion mole numbers adds in ammonium chloride solution, reacts and is precipitated thing after 5 minutes, solid-liquid separation, the zinc sulphide obtaining respectively and ammonium chloride solution; The ammonium chloride solution finally obtaining all can return to step (2) and recycle, and when mol ratio is during higher than 2:1, the mol ratio that adds ammonium chloride to be adjusted to ammonium ion and chlorion is: 2:1.
Embodiment 5:
Identical with embodiment 3 steps (1)～(3), the ammonium chloride leach liquor obtaining in step (3), adopt precipitation agent (ammonium sulfide) directly to process 5 minutes, precipitation agent consumption is zinc, lead, copper, cadmium and silver-colored integral molar quantity 1.5 times in ammonium chloride, obtains lead sulfide zinc ore concentrate and the ammonium chloride solution of cupric silver after filtration.
Embodiment 6:
Identical with embodiment 4 steps (1)～(3), the ammonium chloride leach liquor obtaining in step (3), adopt precipitation agent (sodium sulphite) directly to process 30 minutes, precipitation agent consumption is zinc, lead, copper, cadmium and silver-colored integral molar quantity 0.5 times in ammonium chloride, obtains lead sulfide zinc ore concentrate and the ammonium chloride solution of cupric silver after filtration.
Claims (10)
1. alkali ammonium sulphur coupled method is processed a method for zinc hydrometallurgy waste residue, it is characterized in that concrete steps comprise:
(1) by the roasting 0.5～3 hour at 300～900 ℃ of temperature of the waste residue of zinc hydrometallurgy, making siderotil or zinc ferrite thing inversion of phases in slag is the fired slags of zinc sulfate, lead sulfate and ferric oxide;
(2) by solid-to-liquid ratio 1:4～1:10 g/mL, with sodium hydroxide solution, at 30～160 ℃, leach the fired slags that obtains in step (1) 0.5～3 hour, solid-liquid separation obtains alkali and soaks slag and alkali immersion liquid, and recycle is returned in alkali immersion liquid;
(3) alkali step (2) being obtained soaks slag and immerses 0.5～3h in the mixing solutions that temperature is 90～125 ℃ of ammonium chlorides and ammoniacal liquor, obtains ammonium chloride leach liquor and leached mud after filtration;
(4) zinc powder of 0.5～3.0 times of the total mole number of copper in ammonium chloride leach liquor, lead, cadmium and silver is reduced to 5～60 minutes to reclaim copper, lead, cadmium and the silver in leach liquor at 20～60 ℃, after solid-liquid separation, obtain respectively the cadmia of leaded, copper, cadmium and silver and containing the ammonium chloride solution of zine ion;
(5) by ammonium chloride solution in step (4), the precipitation agent containing 0.5～2.0 times of zine ion mole number adds in ammonium chloride solution, reacts and is precipitated thing after 5～30 minutes, solid-liquid separation, the zinc sulphide obtaining respectively and ammonium chloride solution.
2. alkali ammonium sulphur coupled method is processed a method for zinc hydrometallurgy waste residue, it is characterized in that concrete steps comprise:
(1) by the roasting 0.5～3 hour at 300～900 ℃ of temperature of the waste residue of zinc hydrometallurgy, making siderotil or zinc ferrite thing inversion of phases in slag is the fired slags of zinc sulfate, lead sulfate and ferric oxide;
(2) by solid-to-liquid ratio 1:4～1:10 g/mL, with sodium hydroxide solution, at 30～160 ℃, leach the fired slags that obtains in step (1) 0.5～3 hour, solid-liquid separation obtains alkali and soaks slag and alkali immersion liquid, and recycle is returned in alkali immersion liquid;
(3) alkali step (2) being obtained soaks slag and immerses 0.5～3h in the mixing solutions that temperature is 90～125 ℃ of ammonium chlorides and ammoniacal liquor, obtains ammonium chloride leach liquor and leached mud after filtration;
(4) adopt precipitation agent directly to process ammonium chloride leach liquor 5～30 minutes, precipitation agent consumption is zinc, lead, copper, cadmium and silver-colored integral molar quantity 0.5～1.5 times in ammonium chloride leach liquor, obtains lead sulfide zinc ore concentrate and the ammonium chloride solution of cupric silver after filtration.
3. alkali ammonium sulphur coupled method according to claim 1 and 2 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: the waste residue of described zinc hydrometallurgy is plumbous silver-colored slag or iron vitriol slag, its composition and percentage range: zinc 3～15 wt %, plumbous 3～10 wt%, silver 50～500 g/t, iron 5～30 wt%, all the other are impurity.
4. alkali ammonium sulphur coupled method according to claim 1 and 2 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: the concentration of described sodium hydroxide solution is 30～100 g/L.
5. alkali ammonium sulphur coupled method according to claim 1 and 2 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: during the alkali immersion liquid recycle that obtains in described step (2), concentration is during lower than 30 g/L, recycling while adding sodium hydroxide to adjust concentration to 30～100 g/L.
6. alkali ammonium sulphur coupled method according to claim 1 and 2 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: in described step (3), ammonia concn is 1
~30g/L, the concentration of ammonium chloride solution is 100～400g/L, in mixing solutions, the mol ratio of ammonium ion and chlorion is 1:1～2:1.
7. alkali ammonium sulphur coupled method according to claim 1 and 2 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: in described step (3), alkali soaks in slag immersion ammonium chloride and ammoniacal liquor mixing solutions, and solid-to-liquid ratio is 1:10 g/mL.
8. alkali ammonium sulphur coupled method according to claim 1 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: the granularity of described zinc powder is-100 orders, and purity is more than 99%.
9. alkali ammonium sulphur coupled method according to claim 1 and 2 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: described precipitation agent is ammonium sulfide or sodium sulphite.
10. alkali ammonium sulphur coupled method according to claim 1 and 2 is processed the method for zinc hydrometallurgy waste residue, it is characterized in that: the described ammonium chloride solution finally obtaining all returns to step (3) and recycles, when the mol ratio of ammonium ion and chlorion is during lower than 1:1, the mol ratio that adds ammoniacal liquor to be adjusted to ammonium ion and chlorion is: 2:1～1:1; When mol ratio is during higher than 2:1, the mol ratio that adds ammonium chloride to be adjusted to ammonium ion and chlorion is: 2:1～1:1.
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| CN109970105B (en) * | 2019-04-24 | 2021-04-16 | 西南科技大学 | Method for cleanly recovering iron in zinc hydrometallurgy process |
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| CN112176199B (en) * | 2020-09-27 | 2022-08-02 | 衡阳市大宇锌业有限公司 | Process for efficiently purifying zinc oxide from zinc smelting waste residues |
| CN112981121B (en) * | 2021-02-05 | 2022-08-23 | 江西思远再生资源有限公司 | Method for treating cadmium-containing waste residue and preparing high-purity metal cadmium by precipitation method |
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